Method for thiosulfate leaching of precious metal-containing materials

ABSTRACT

Processes are provided for recovering precious metals from refractory materials using thiosulfate lixiviants. The processes can employ lixiviants that include at most only small amounts of copper and/or ammonia and operate at a relatively low pH, reduction of polythionates, inert atmospheres to control polythionate production, and electrolytic solutions which provide relatively high rates of precious metal recovery.

CROSS REFERENCE TO RELATED APPLICATION

The present application is a divisional of U.S. application Ser. No.13/020,505, filed Feb. 3, 2011, which is a divisional of U.S.application Ser. No. 12/700,525, filed Feb. 4, 2010, which is acontinuation of U.S. application Ser. No. 10/836,480, filed Apr. 30,2004, now U.S. Pat. No. 7,704,298, which is a divisional application ofU.S. application Ser. No. 10/446,548, filed May 27, 2003, now U.S. Pat.No. 7,066,983, which is a divisional application of U.S. applicationSer. No. 09/852,699, filed May 11, 2001, now U.S. Pat. No. 6,660,059,which claims priority from U.S. Provisional Application Ser. No.60/205,472, filed May 19, 2000, all of which are incorporated herein bythis reference.

FIELD OF THE INVENTION

The present invention is directed generally to the recovery of preciousmetals from precious metal-containing material and specifically to therecovery of precious metals from precious metal-containing materialusing thiosulfate lixiviants.

BACKGROUND OF THE INVENTION

A traditional technique for recovering precious metal(s) from preciousmetal-containing ore is by leaching the material with a cyanidelixiviant. As used herein, a “precious metal” refers to gold, silver,and the platinum group metals (e.g., platinum, palladium, ruthenium,rhodium, osmium, and iridium). Many countries are placing severelimitations on the use of cyanide due to the deleterious effects ofcyanide on the environment. Incidents of fish and other wildlife havingbeen killed by the leakage of cyanide into waterways have been reported.The limitations being placed on cyanide use have increased substantiallythe cost of extracting precious metal(s) from ore, thereby decreasingprecious metal reserves in many countries. Cyanide is also unable torecover precious metals such as gold from refractory ores without apretreatment step. “Refractory ores” refer to those ores that do notrespond well to conventional cyanide leaching. Examples of refractoryores include sulfidic ores (where at least some of the precious metalsare locked up in the sulfide matrix), carbonaceous ores (where theprecious metal complex dissolved in the lixiviant adsorbs ontocarbonaceous matter in the ores), and sulfidic and carbonaceous ores.

Thiosulfate has been actively considered as a replacement for cyanide.Thiosulfate is relatively inexpensive and is far less harmful to theenvironment than cyanide. Thiosulfate has also been shown to beeffective in recovering precious metals from pretreated refractorypreg-robbing carbonaceous ores and sulfidic ores. As used herein,“preg-robbing” is any material that interacts with (e.g., adsorbs orbinds) precious metals after dissolution by a lixiviant, therebyinterfering with precious metal extraction, and “carbonaceous material”is any material that includes one or more carbon-containing compounds,such as humic acid, graphite, bitumins and asphaltic compounds.

Where gold is the precious metal, thiosulfate leaching techniques havetypically relied on the use of copper ions to catalyze and acceleratethe oxidation of gold, ammonia to facilitate the formation andstabilization of cupric ammine ions and/or a pH at pH 9 or above tomaintain a region of stability where both the cupric ammine and goldthiosulfate complexes are stable.

It is well known in the art that the catalytic effect of copper andammonia in conventional thiosulfate leaching of gold is described by thefollowing sequence of reactions.

Formation of the cupric ammine complex:

Cu²⁺+4NH₃→Cu(NH₃)₄ ²⁺  (1)

Oxidation of gold by cupric ammine, gold complexation as thegold-thiosulfate anion, and reduction of the cupric ammine to cuprousthiosulfate:

Au+Cu(NH₃)₄ ²⁺+5S₂O₃ ²⁻→Au(S₂O₃)₂ ³⁻+Cu(S₂O₃)₃ ⁵⁻+4NH₃  (2)

Oxidation of the cuprous thiosulfate back to cupric ammine with oxygen:

Cu(S₂O₃)₃ ⁵⁻+4NH₃+¼O₂+½H₂O→Cu(NH₃)₄ ²⁺+3S₂O₃ ²⁻+OH⁻  (3)

Summing equations (2) and (3) yields the overall thiosulfate leachreaction for gold:

Au+2S₂O₃ ²⁻+¼O₂+H₂O→Au(S₂O₃)₂ ³⁻+OH⁻  (4)

It can be seen from the above equations that copper and ammonia act ascatalysts in that they are neither produced nor consumed in the overallleach reaction.

Cupper and ammonia can be a source of problems. Added copper tends toprecipitate as cupric sulfide, which is speculated to form a passivelayer on gold, thereby inhibiting gold leaching as well as increasingcopper and thiosulfate consumption:

Cu²+S₂O₃ ²⁻+2OH⁻→CuS+SO₄ ²⁻+H₂O  (5)

Rapid oxidation of thiosulfate by cupric ammine also occurs, leading toexcessive degradation and loss of thiosulfate:

2Cu(NH₃)₄ ²⁺+8S₂O₃ ²⁻→2Cu(S₂O₃)₃ ⁵⁻+S₄O₆ ²⁻+8NH₃  (6)

Loss of ammonia by volatilization occurs readily, particularly inunsealed gas-sparged reactors operating at pH greater than 9.2, leadingto excessive ammonia consumption:

NH₄ ⁺+OH⁻→NH_(3(aq))+H₂O→NH_(3(g))+H₂O  (7)

Like cyanide, copper and ammonia are highly toxic to many aquaticlifeforms and are environmentally controlled substances.

Other problems encountered with thiosulfate leaching include difficultyin recovering gold out of solution as a result of the formation ofpolythionates, such as tetrathionate and trithionate, which adsorbcompetitively with gold onto adsorbents, such as resins. The formationof polythionates further increases thiosulfate consumption per unit massof processed ore.

SUMMARY OF THE INVENTION

These and other needs have been addressed by the methodologies andsystems of the present invention. The methodologies can recover preciousmetals from a variety of materials, including refractory carbonaceous orsulfidic ores, double refractory ores (e.g., ores containing bothsulfide-locked gold and carbonaceous preg-robbing matter), oxide ores,nonrefractory sulfidic ores, and ores also containing copper mineralsand other materials derived from such ores (e.g., concentrates,tailings, etc.).

In one embodiment, a thiosulfate leaching process is provided thatincludes one or more of the following operating parameters:

(a) an oxygen partial pressure that is preferably superatmospheric andmore preferably ranges from about 4 to about 500 psia;

(b) a leach slurry pH that is preferably less than pH 9;

(c) a leach slurry that is preferably at least substantially free of(added) ammonia and more preferably contains less than 0.05M (added)ammonia such that the leach slurry has a maximum total concentration ofammonia of preferably less than 0.05M and more preferably no more thanabout 0.025M;

(d) a leach slurry that is preferably at least substantially free of(added) copper ion and more preferably contains no more than about 15ppm (added) copper ions;

(e) an (added) sulfite concentration that is preferably no more thanabout 0.01M such that the slurry has a maximum total concentration ofsulfite of preferably no more than about 0.02M and more preferably nomore than about 0.01M; and/or

(f) a leach slurry temperature preferably ranging from about 20 to about100° C. and more preferably from about 20 to about 80° C.

The foregoing parameters can yield a high level of precious metalextraction from the precious metal-containing material, which can be atleast about 70% and sometimes at least about 80%.

The thiosulfate lixiviant can be derived from any suitable form(s) ofthiosulfate, such as sodium thiosulfate, calcium thiosulfate, potassiumthiosulfate and/or ammonium thiosulfate. Sodium and/or calciumthiosulfate are preferred.

The leaching process can be conducted by any suitable technique. Forexample, the leaching can be conducted in situ, in a heap or in an openor sealed vessel. It is particularly preferred that the leaching beconducted in an agitated, multi-compartment reactor such as anautoclave.

The precious metal can be recovered from the pregnant leach solution byany suitable technique. By way of example, the precious metal can berecovered by resin adsorbtion methods such as resin-in-pulp,resin-in-solution, and resin-in-leach or by solvent extraction,cementation, electrolysis, precipitation, and/or combinations of two ormore of these techniques.

Reducing or eliminating the need to have copper ions and/or ammoniapresent in the leach as practiced in the present invention can providesignificant multiple benefits. First, the cost of having to add copperand ammonia reagents to the process can be reduced significantly oreliminated. Second, environmental concerns relating to the presence ofpotentially harmful amounts of copper and ammonia in the tailings orother waste streams generated by the process can be mitigated. Third,the near-absence or complete absence of copper and ammonia in the leachcan provide for a much more reliable and robust leaching process,yielding more stable leachates, able to operate over a wider pH andoxidation-reduction potential (ORP) range than is possible withconventional thiosulfate leaching. The latter process must operate inthe relatively narrow window of pH and ORP where both the cupric amminecomplex and the gold thiosulfate complex co-exist. With the process ofthe present invention, the pH of the thiosulfate lixiviant solution inthe leaching step can be less than pH 9 and the ORP less than 200 mV(referenced to the standard hydrogen electrode). Fourth, minimizing theamount of copper in the system can lead to increased loading of goldonto resins due to reduced competitive adsorption of copper ions. Resinelutions are also simplified as little, if any copper, is on the resin.Finally, the near-absence or complete absence of copper and ammonia inthe leach can reduce or eliminate entirely a host of deleterious sidereactions that consume thiosulfate and are otherwise difficult orimpossible to prevent.

The elimination or near elimination of sulfite from the thiosulfateleach also can have advantages. Sulfite can depress the rate ofdissolution of precious metal from the precious metal-containingmaterial by reducing significantly the oxidation reduction potential(ORP) of the leach solution or lixiviant. As will be appreciated, therate of oxidation of the gold (and therefore the rate of dissolution ofthe gold) is directly dependent on the ORP.

In another embodiment, an extraction agent is preferably contacted witha pregnant (precious metal-containing) thiosulfate leach solution at atemperature of less than about 70° C. and more preferably less thanabout 60° C. in the substantial absence of dissolved molecular oxygen toisolate the precious metal and convert polythionates in the pregnantleach solution into thiosulfate. In one configuration, the extractionagent is an adsorbent, such as a resin, which loads the precious metalonto the adsorbent. As used herein, an “adsorbent” is a substance whichhas the ability to hold molecules or atoms of other substances on itssurface. Examples of suitable resin adsorbents include weak and strongbase resins such as “DOWEX 21K”, manufactured by Dow Chemical. Inanother configuration, the extraction agent is a solvent extractionreagent that extracts the precious metals into an organic phase, fromwhich the precious metals can be later recovered. As will beappreciated, the detrimental polythionates decompose into thiosulfate inthe substantial absence of dissolved molecular oxygen.

In yet another embodiment, the pregnant leach solution from athiosulfate leaching step is contacted, after the leaching step, with areagent to convert at least about 50% and typically at least most ofpolythionates (particularly trithionate and tetrathionate) intothiosulfate. The reagent or reductant can be any suitable reactant toconvert polythionates into thiosulfate, with any sulfide, and/orpolysulfide (i.e., a compound containing one or a mixture of polymericion(s) S_(x) ²⁻, where x=2-6, such as disulfide, trisulfide,tetrasulfide, pentasulfide and hexasulfide) being particularlypreferred. A sulfite reagent can also be used but is generally effectiveonly in converting polythionates of the form S_(x)O₆ ²⁻, where x=4 to 6,to thiosulfate. The sulfite, sulfide, and/or polysulfide can becompounded with any cation, with Groups IA and HA elements of thePeriodic Table, ammonium, and hydrogen being preferred.

In yet another embodiment, a precious metal solubilized in a solution,such as a pregnant leach solution or eluate, is electrowon in thepresence of sulfite. In the presence of sulfite, the precious metal isreduced to the elemental state at the cathode while the sulfite isoxidized to sulfate at the anode. Sulfite is also believed to improvethe precious metal loading capacity of the resin by converting loadedtetrathionate to trithionate and thiosulfate.

In yet another embodiment, the formation of polythionates is controlledby maintaining a (pregnant or barren) thiosulfate leach solution in anonoxidizing (or at least substantially nonoxidizing) atmosphere and/orsparging a nonoxidizing (or at least substantially nonoxidizing) gasthrough the leach solution. As will be appreciated, the atmosphere orgas may contain one or more reductants, such as hydrogen sulfide and/orsulfur dioxide. The molecular oxygen concentration in the atmosphereand/or sparge gas is preferably insufficient to cause a dissolvedmolecular oxygen concentration in the leach solution of more than about1 ppm and preferably of more than about 0.2 ppm. Preferably, the inertatmosphere (or sparge gas) is at least substantially free of molecularoxygen and includes at least about 85 vol. % of any inert gas such asmolecular nitrogen and/or argon. By controlling the amount of oxidant(s)(other than thiosulfate and polythionates) in the atmosphere and/or(pregnant or barren) leach solution the rate or degree of oxidation ofthiosulfates to form polythionates can be controlled.

As used herein, “at least one . . . and”, “at least one . . . or”, “oneor more of . . . and”, “one or more of . . . or”, and “and/or” areopen-ended expressions that are both conjunctive and disjunctive inoperation. For example, each of the expressions “at least one of A, Band C”, “at least one of A, B, or C”, “one or more of A, B, and C”, “oneor more of A, B, or C” and “A, B, and/or C” means A alone, B alone, Calone, A and B together, A and C together, B and C together, and A, Band C together.

BRIEF DESCRIPTION OF THE DRAWINGS

FIG. 1 is a flow schematic of a first embodiment of the presentinvention;

FIG. 2 is a flow schematic of second embodiment of the presentinvention;

FIG. 3 is a flow schematic of a third embodiment of the presentinvention;

FIG. 4 is a flow schematic of a fourth embodiment of the presentinvention;

FIG. 5 is a plot of gold extraction in percent (vertical axis) versusleach time in hours (horizontal axis);

FIG. 6 is another plot of gold extraction in percent (vertical axis)versus leach time in hours (horizontal axis);

FIG. 7 is another plot of gold extraction in percent (vertical axis)versus leach time in hours (horizontal axis);

FIG. 8 is another plot of gold extraction in percent (vertical axis)versus leach time in hours (horizontal axis); and

FIG. 9 is a plot of gold extraction in percent (left vertical axis) andthiosulfate remaining in percent (right vertical axis) versus leach timein hours (horizontal axis);

DETAILED DESCRIPTION

The present invention provides an improved thiosulfate leaching processfor the recovery of precious metals from precious metal-bearingmaterial. The precious metal(s) can be associated with nonpreciousmetals, such as base metals, e.g., copper, nickel, and cobalt. Theprecious metal-bearing material includes ore, concentrates, tailings,recycled industrial matter, spoil, or waste and mixtures thereof. Theinvention is particularly effective for recovering precious metals,particularly gold, from refractory carbonaceous material.

FIG. 1 is a flow chart according to a first embodiment of the presentinvention. The process of the flow chart is particularly effective inrecovering gold from carbonaceous material and oxide material andmixtures thereof.

Referring to FIG. 1, a precious metal-bearing material 100 is subjectedto the steps of wet and/or dry crushing 104 and wet and/or dry grinding108 to reduce the particle size of the material sufficiently to enablethe solids to be suspended in an agitated vessel and to allow for theefficient leaching of the precious metals. Preferably, wet grinding isemployed with the recycled thiosulfate leach solution and water beingused as the liquid component in the slurry. In that event, the slurry112 containing the comminuted material typically contains from about0.05 to about 0.1 M thiosulfates and from about 0.0005 to about 0.025 Mpolythionates. The fully comminuted material particle size is preferablyat least smaller than 80% passing about 48 mesh (300 microns), morepreferably 80% passing about 100 mesh (150 microns), and most preferably80% passing about 200 mesh (75 microns). The typical solids content ofthe slurry 112 ranges from about 20 to about 30 wt. %. As will beappreciated, other techniques can be used to comminute the material tothe desired particle size(s). By way of illustration, blasting can beused alone with or without crushing and grinding and crushing andgrinding can be used alone with or without another comminutiontechnique.

The ground slurry 112 is then thickened 116 to adjust the pulp densityto a value suitable for leaching. The ideal leach pulp density will varyaccording to the type of material being leached. Typically, the pulpdensity ranges from about 20 to about 50% solids by weight, but could beas low as about 1% or as high as about 60%. Thickening 116 willgenerally not be required if the desired pulp density (after wetcomminution or formation of the comminuted material into a slurry) isless than about 20%.

The thickener overflow solution 120 is recycled back to grinding 108 inthe event that wet grinding is employed. Otherwise, the overflowsolution 120 is returned to the optional slurry formation step (notshown).

Fresh makeup thiosulfate is added, as necessary, at any suitablelocation(s), such as to the slurried material during comminution 108and/or in the thickener 116, to the underflow or overflow solution 124,120, to leaching 132 and/or to the regenerated thiosulfate solution 128(discussed below). In any event, the optimum solution thiosulfateconcentration to maintain during leaching 132 will depend on the natureof the material being leached, but will preferably range from about0.005 to about 2 molar (M), more preferably about 0.02 to about 0.5 M,and even more preferably from about 0.05 to about 0.2 M. The source ofmakeup thiosulfate can be any available thiosulfate-containing compound,such as sodium thiosulfate, potassium thiosulfate, calcium thiosulfate,or any other thiosulfate-containing material or thiosulfate precursor.Ammonium thiosulfate can also be used but its use is less preferred forenvironmental reasons. Alternatively, thiosulfate can be generated insitu or in a separate step by reaction of elemental sulfur with a sourceof hydroxyl ions, in accordance with the following reaction:

2(x+1)S+60H⁻→S₂O₃ ²⁻+2S_(x) ²⁻+3H₂O  (8)

where x=3-6, or by reaction of bisulfide with bisulfite:

2HS⁻+4HSO₃ ⁻→3S₂O₃ ²⁻+3H₂O  (9)

or by reaction of elemental sulfur with sulfite:

S+SO₃ ²⁻→S₂O₃ ²⁻  (10)

If the desirable temperature is above ambient, it may be desirable torecover waste heat for recycle to leaching. In that event, the underflowslurry 124 is directed through an indirect heat exchanger 136,preferably a shell and tube heat exchanger system in which the hotslurry from resin-in-pulp pretreatment 140 (discussed below) is passedthrough the inner tubes and the cold feed (or underflow) slurry 140 ispassed through the annular space between the tubes (or vice versa). Inthis way waste heat is transferred from the leached slurry 144 to thefeed (or underflow) slurry 124, reducing the amount of new heat thatmust be added in leaching 132 to maintain the desired leach temperature.Typically, the approach temperature of the incoming feed slurry 148 isfrom about 2 to about 5° C. below the leach temperature (discussedbelow) and heat is added to the leach vessel by suitable techniques tomakeup the difference.

The heated slurry 148 is subjected to leaching 132 in the presence ofoxygen and thiosulfate. Leaching is conducted in the presence of anoxygen-enriched atmosphere at atmospheric pressure, or at a pressureabove atmospheric pressure using an oxygen-containing gas to reduce oreliminate the need for the presence of copper and/or ammonia in theleach. Using gold as an example, the thiosulfate leaching of preciousmetal-bearing material in the absence or substantial absence of copperand ammonia under elevated oxygen partial pressure can be illustrated bythe following reaction:

$\begin{matrix}{{{Au} + {2S_{2}O_{3}^{2^{-}}} + {\frac{1}{4}O_{2}} + {{1/2}\; H_{2}O}}->{{{Au}\left( {S_{2}O_{3}} \right)}_{2}^{3^{-}} + {OH}^{-}}} & (11)\end{matrix}$

The increased oxygen partial pressure in the leach increases the rate ofthe above reaction in the absence or near absence of copper and ammonia.To accomplish this goal, the oxygen-containing gas may includeatmospheric air, or it may include relatively pure (95%+) oxygen such asthat produced from any commercially available oxygen plant, or it mayinclude any other available source of oxygen. The desired oxygen partialpressure (PO₂) maintained during leaching will depend on the materialbeing leached, but it will be at least higher than that provided undernormal ambient conditions by air at the elevation the process isapplied. Thus, if the process is practiced at sea level for example theoxygen partial pressure will be in excess of about 3 pounds per squareinch absolute pressure (psia) to as high as about 500 psia, preferablyfrom about 10 to about 115 psia, and most preferably from about 15 toabout 65 psia. The total operating pressure is the sum of the molecularoxygen partial pressure and the water vapor pressure at the temperatureemployed in the leaching step 132, or preferably ranges from about 15 toabout 600 psia and more preferably from about 15 to about 130 psia.

The leaching temperature will be dictated by the type of material beingleached. The temperature will vary typically from about 5° C. to about150° C., preferably from about 20 to about 100° C., and most preferablyfrom about 40 to about 80° C. Higher temperatures accelerate theleaching of precious metals but also accelerate the degradation ofthiosulfate. If required, a source of makeup heat such as steam is addedto the leach reactors to maintain the desired temperature.

The leaching retention time is dependent on the material being leachedand the temperature, and will range from about 1 hour to 96 hours,preferably from about 2 to about 16 hours, and most preferably fromabout 4 to about 8 hours.

The absence or substantial absence of copper and/or ammonia in the leachgreatly simplifies the process. Elimination or near-elimination ofammonia and copper from the leach provides the advantage of allowing fora consistently high and reproducible precious metal extraction over abroader pH range than was previously possible with the other thiosulfateleaching processes. Preferably, the (added and/or total solution) copperconcentration is no more than about 20 ppm, more preferably no more thanabout 15 ppm, and even more preferably no more than about 10 ppm whilethe (added and/or total solution) ammonia concentration is no more thanabout 0.05 M, more preferably no more than about 0.03 M, and even morepreferably no more than about 0.01 M. In the present invention leachingcan be operated at about pH 7-12, preferably about pH 8-11, morepreferably about pH 8-10, and even more preferably at a pH less than pH9. The oxidation-reduction potential (ORP) preferably ranges from about100 to about 350 mV and more preferably from about 150 to about 300 mV(vs. the standard hydrogen electrode (SHE)).

Oxidative degradation of thiosulfate ultimately to sulfate can alsooccur, possibly by the following sequence of reactions that involve theformation of intermediate polythionates (polythionates can berepresented by S_(n)O₆ ²⁻, where n=2-6):

Tetrathionate formation: 2S₂O₃ ²⁻+½O₂+H₂O→S₄O₆ ²⁻+2OH⁻  (12)

$\begin{matrix}{{{Trithionate}\mspace{14mu} {formation}\text{:}}{{{3S_{4}O_{6}^{2^{-}}} + {\frac{5}{2}O_{2}} + {H_{2}O}}->{{4S_{3}O_{6}^{2^{-}}} + {2H^{+}}}}} & (13)\end{matrix}$Sulfite formation: S₃O₆ ²⁻+½O₂+2H₂O→3SO₃ ²⁻+4H⁺  (14)

Sulfate formation: 2SO₃ ²⁻+O₂→2SO₄ ²⁻  (15)

Overall: S₂O₃ ²⁻+2O₂+H₂O→2SO₄ ²⁻+2H⁺  (16)

Oxidative degradation of thiosulfate to polythionates and sulfates isaccelerated markedly in the presence of copper ions and/or ammonia. Theoxidative degradation reactions are slowed considerably at elevatedoxygen partial pressure in the absence or near-absence of copper andammonia.

The leaching step 132 may be conducted in a batch or continuous basisbut continuous operation is preferred. Continuous leaching is carriedout in a multiple series of one or more reactors that are agitatedsufficiently to maintain the solids in suspension. Agitation may beaccomplished by mechanical, pneumatic or other means. In a preferredconfiguration, gassing impellers are employed, such as those disclosedin U.S. Pat. No. 6,183,706 and copending U.S. patent application Ser.No. 09/561,256, filed Apr. 27, 2000, which are incorporated herein byreference. Such impellers can significantly enhance the amount ofdissolved molecular oxygen in the leach solution. Leaching may also becarried out in a multi-compartment autoclave containing one or morecompartments, (with 4 to 6 compartments being preferred) similar indesign to the autoclaves used to pressure oxidize sulfide-bearing oresor concentrates. However, owing to the non-acidic, moderate temperature,relatively mild conditions employed in the present invention, theautoclave materials of construction are much less expensive than thosefound to be necessary when oxidizing sulfide minerals. The latterautoclaves are normally constructed of a steel shell fitted with a leadliner and refractory brick liner and containing metallic componentsconstructed of titanium or other expensive corrosion-resistant alloys.The leach reactors and contained metallic components employed by thepresent invention can be simply constructed of stainless steel and donot require lead or brick liners.

The extraction of precious metals in the leaching step 132 is relativelyhigh, particularly for carbonaceous ores. Typically, at least about 50%,more typically at least about 70%, and even more typically at leastabout 80% of the precious metal in the precious metal-containingmaterial is extracted or solubilized into the pregnant leach solution144. The concentration of the dissolved precious metal in the pregnantleach solution typically ranges from about 0.05 to about 100 ppm andmore typically from about 1 to about 50 ppm.

The pregnant leach slurry 144 containing the precious metal-bearingleach solution and gold-depleted solid residue may optionally bedirected to RIP pretreatment 140 to reduce the concentration ofpolythionates in solution. As will be appreciated, the molecular oxygensparged through the leach slurry in the leaching step 132 will oxidize aminor portion of the thiosulfate into polythionates. Polythionates havethe undesired effect of reducing the loading of precious metals on toresin by competitive adsorption. Lowering the polythionate concentrationwill have the beneficial effect of increasing the loading of preciousmetals on to resin, thereby improving the efficiency of resin recoveryof precious metals.

The RIP pretreatment step 140 can be performed using any one or more ofa number of techniques for converting polythionates to other compoundsthat do not compete with the precious metal for collection by theextraction agent.

In one embodiment, a polythionate reductant is added to the slurry 144to reduce polythionates to thiosulfates. Any of a number of reductantsare suitable for performing the conversion.

By way of example, a sulfide-containing reagent can reduce thepolythionates back to thiosulfate, as shown by the following reactions:

2S₄O₆ ²⁻+S²⁻+3/2H₂O→9/2S₂O₃ ²⁻+3H⁺  (17)

S₃O₆ ²⁻+S²⁻→2S₂O₃ ²⁻  (18)

Any reagent that releases sulfide ions on dissolution will suffice, suchas sodium bisulfide, NaHS, sodium sulfide, Na₂S, hydrogen sulfide gas,H₂S, or a polysulfide. The use of a sulfide reagent has the advantagesof rapidly and efficiently reducing polythionates to thiosulfate atambient or moderately elevated temperature. The treatment can be carriedout in an agitated reactor in batch mode or in a series of 1-4 reactorsoperating in continuous mode, or in a multi-compartment autoclave.Alternatively the treatment can be carried out in a pipe reactor orsimply by injecting sulfide ions in the piping system directing theleach slurry to gold recovery, or the first stage of RIP. The treatmentis carried out at a controlled pH of about pH 5.5 to about pH 10.5,preferably about pH 7 to about pH 10, most preferably less than about pH9. The temperature employed can range from about 20° C. to about 150°C., preferably from about 40 to about 100° C., more preferably fromabout 40 to about 80° C., and even more preferably from about 60 toabout 80° C. The retention time can range from as low as 5 minutes,preferably greater than 30 minutes, most preferably from about 1 toabout 3 hours.

Alternatively, a sulfite-containing reagent can also reducepolythionates to thiosulfates as shown by the following reaction:

S₄O₆ ²⁻+SO₃ ²⁻→S₂O₃ ²⁻+S₃O₆ ²⁻  (19)

Sulfite treatment is effective in reducing tetrathionate quickly, but adisadvantage is it is ineffective in reducing trithionate. The sulfitecan be added in any form and/or can be formed in situ. For example,sulfite can be added in the form of sodium metabisulfite or sulfurdioxide.

When using sulfite, the temperature of the leach slurry in the RIPpretreatment 140 is preferably less than 60° C. to inhibit, at leastsubstantially, the precipitation of precious metal(s) from the leachslurry 144. More preferably, the RIP pretreatment 140 with sulfite isperformed at a temperature in the range of from about 10 to about 50° C.and even more preferably at ambient temperature.

When using sulfite, the residence time of the leach slurry 144 in theregeneration step 140 is preferably at least about 1 second, morepreferably greater than about 5 minutes, and even more preferablygreater than about 10 minutes and no more than about 1 hour, with about15-30 minutes being most preferable.

The pH of the leach slurry during sulfite treatment typically rangesfrom about pH 5.5 to about pH 10.5 and more typically from about pH 7 toabout pH 10.

Other suitable polythionate reductants include hydrogen, fine, reactiveelemental sulfur, carbon monoxide, and mixtures thereof.

In another embodiment, the pretreatment step 140 is performed bymaintaining the temperature of the leach slurry at a sufficiently highvalue in the absence of oxygen to effect the following hydrolyticdisproportionation reactions:

4S₄O₆ ²⁻+5H₂O→7S₂O₃ ²⁻+2SO₄ ²⁻+10H⁺  (20)

S₃O₆ ²⁻+H₂O→S₂O₃ ²⁻+SO₄ ²⁻+2H⁺  (21)

Hydrolytic treatment can be carried out in an agitated reactor in batchmode or in a series of 1-4 reactors operating in continuous mode, or amulti-compartment autoclave. The temperature is preferably maintained inthe range of from about 60 to about 150° C., preferably of from about 70to about 100° C., and most preferably of from about 80 to about 90° C.by adding a source of heat, such as steam. The retention time typicallyranges from about 15 minutes to 8 hours, preferably from about 1 toabout 6 hours, and most preferably from about 2 to about 4 hours.Hydrolytic treatment is generally less preferable than sulfide treatmentbecause the former method results in irreversible loss of some of thepolythionate to sulfate.

Alternatively, any or all of the above-techniques for convertingpolythionate(s) into thiosulfate can be combined in the same processconfiguration.

In a preferred embodiment, the reductant used to convert polythionatesinto thiosulfates is the sulfide reagent. Sulfide addition is preferredbecause one sulfide reacts with one tri- or two tetrathionates to formmultiple thiosulfates without any sulfur-containing byproducts. Sulfiteaddition only reduces tetrathionate and is not capable of reducingtrithionate at common operating temperatures and pH's. Heating of theleach solution is energy intensive and produces byproducts. Trithionateand tetrathionate are each converted into thiosulfate, sulfate, andhydrogen ions, thus the thiosulfate yield is not as high as with sulfideaddition.

RIP pretreatment 140 can be performed in any suitable vessel(s),preferably agitated. Preferably, RIP pretreatment is performed in aseries of tanks or in a multistaged vessel.

The addition of a sulfide such as NaHS is preferred. Preferably, theamount of the reductant generally and, sulfide reagent specificallyadded to the slurry 144 is sufficient to convert at least most of thepolythionates into thiosulfate. The amount of sulfide contacted with theslurry 144 preferably is at least about 100 to about 150% of thestoichiometric amount required to convert at least substantially all ofthe polythionates in the slurry into thiosulfates. Typically, at leastabout 50%, more typically at least most, and even more typically fromabout 80 to about 95% of the polythionates are converted intothiosulfates in RIP pretreatment 140.

The temperature of the slurry 144 preferably is at least about 60° C.and the ORP of the exiting slurry 152 is at least below about 100 mV(SHE) and more preferably ranges from about −100 to about 100 mV (SHE)to substantially minimize precious metal precipitation.

The exiting RIP pretreated slurry 152 is passed through heat exchanger136 and conditioned in a conditioner 156 to resolubilize any preciousmetal precipitated during RIP pretreatment 140 and/or heat exchange 136.Conditioning 156 is performed in an agitated single- ormulti-compartment vessel which has an oxidizing atmosphere, such as air,to cause solubilization of the precious metal precipitates. Althoughpolythionates will form in the presence of an oxidant, such as molecularoxygen, the rate of conversion of thiosulfate to polythionates is muchslower than the rate of precious metal solubilization. Preferably, theresidence time (at ambient temperature and pressure) is selected suchthat at least about 95% of the precious metal precipitates aresolubilized while no more than about 5% of the thiosulfate is convertedinto polythionates. Preferably, the slurry residence time inconditioning 156 is no more than about 12 hrs and more preferably rangesfrom about 1 to about 6 hrs,

The conditioned slurry 160 is next subjected to resin-in-pulp treatment164 to extract the precious metal from the conditioned slurry 160. Theresin-in-pulp step 164 can be performed by any suitable technique withany suitable ion exchange resin. Examples of suitable techniques includethat discussed in U.S. patent application Ser. No. 09/542,736, filed inJune, 2000, entitled “A Process for Recovering Gold from ThiosulfateLeach Solutions and Slurries with Ion Exchange Resins”, to Thomas, etal.; U U.S. patent application Ser. No. 09/034,846, filed Mar. 4, 1998,entitled “Method for Recovering Gold from Refractory Carbonaceous Ores”;and U.S. Pat. Nos. 5,536,297 and 5,785,736, all of which areincorporated herein by reference. Preferred resins include anionexchange resins, preferably a strong base resin including a quaternaryamine attached to a polymer backbone. A strong base resin is preferredover a weak base resin. The precious metal loading capacity of a strongbase resin is typically greater than that of a weak base resin, suchthat a lower volume of resin is required. Gel resins and macroporousresins are suitable. Suitable resins include all commercial strong-baseresins of either Type I (triethylamine functional groups) or Type II(triethyl ethanolamine functional groups). Specific strong-base ionexchange resins include “A500” manufactured by Purolite, “A600”manufactured by Purolite, “21K” manufactured by Dow Chemical, “AmberliteIRA 410” manufactured by Rohm and Haas, “Amberlite IRA 900” manufacturedby Rohm and Haas, and “Vitrokele 911” supplied by Signet. Because theRIP pretreatment and resin-in-pulp steps 140 and 164 are preferablyperformed in the same vessel (though they may be performed in differentvessels), the temperature, leach slurry pH, and residence time typicallydepend on which of the above techniques are used to reduce thepolythionate concentration.

Resin-in-pulp treatment can be performed in any suitable vessel. Apreferred vessel is a Pachuca tank, which is an air-agitated, conicalbottomed vessel, with air being injected at the bottom of the cone. Anadvantage of the Pachuca system is reduced resin bead breakage andimproved dispersion of the resin beads in the slurry as compared tomechanically agitated systems. The RIP recovery is preferably carriedout in four or more tanks connected in series, more preferably betweenfour and eight such Pachuca tanks.

During resin-in-pulp 164, the resin will become “loaded” with thedissolved precious metals. Typically, at least about 99% and moretypically at least about 99.8% of the precious metal(s) in the leachslurry will be “loaded” or adsorbed onto the resin.

To inhibit the formation of polythionates and the consequent preciousmetal recovery problems and increased reagent consumption, the leachslurry can be maintained in an inert (or an at least substantiallynonoxidizing) atmosphere and/or an inert (or an at least substantiallynonoxidizing) gas can be sparged through the leach slurry. Theatmosphere is preferably maintained (and/or gas sparging used) duringRIP pretreatment 140 and resin-in-pulp 164. As used herein, “inert”refers to any gas which is at least substantially free of oxidants, suchas molecular oxygen, that can cause thiosulfate to be converted into apolythionate. For example, an “inert” gas would include a reducing gas.Typically, the inert atmosphere will include at least about 85 vol % ofan inert gas, preferably nitrogen gas, and no more than about 5 vol %oxidants, such as oxygen gas, that can cause thiosulfate conversion intoa polythionate. The molecular nitrogen can be a byproduct of the oxygenplant that is employed in the leaching step to provide superatmosphericpartial pressures of oxygen gas. As will be appreciated, the leachslurry 144 during transportation between the leaching and RIPpretreatment steps 132 and 140 and if applicable from the RIPpretreatment and resin-in-pulp steps 140 and 164 (except duringconditioning 156) is typically in a conduit that is not open to thesurrounding atmosphere. If the leach slurry is open to an atmosphereduring transportation in either or both of these stages, the leachslurry should be maintained in the presence of the inert atmosphereduring any such transportation.

While not wishing to be bound, it is believed that sparging is moreeffective than an inert atmosphere without sparging in controllingpolythionate production. Sparging appears to inhibit molecular oxygeningress into the solution, even where the reactor is open to the ambientatmosphere, because of the outflow of inert gas from the surface of thesolution.

The barren leach slurry 168 (which will typically contain no more thanabout 0.01 ppm precious metals or 1% of the precious metal(s) in theleach solution 144) is subjected to one or more stages of countercurrent decantation (“CCD”) 172. In CCD 172, the solids are separated inthe underflow 176 from the barren leach (or overflow) solution 180 andsent to the tailings pond. The barren leach solution 180 is separated inthe overflow from the solids and forwarded to regeneration step 184 toconvert polythionates to thiosulfate. As will be appreciated, CCDperforms liquid/solid separation, provides water balancing in thecircuit, and prevents build up of impurities in the leach circuit byremoving a portion of the leach solution with the solids.

Regeneration 184 can be performed in one or more vessel(s) and/or by inline sulfide (and/or sulfite) addition to a conduit carrying thestripped lixiviant solution. If a number of the techniques are employed,they can be performed simultaneously (in the same reactors) orsequentially (in different reactors), as desired.

The regenerated lixiviant solution 128 is recycled to the grinding step108 along with the thickener overflow 120 and ultimately to the leachingstep 132.

The loaded resin 188 is screened 190 and washed with water to remove anyleach slurry (liquid and/or leached material) from the resin beads.

The recovered beads 192 are contacted with an eluant to strip or elute194 adsorbed precious metal into the eluate and form a pregnant solution196 containing typically at least most (and more typically at leastabout 95%) of the precious metal on the resin and a stripped resin 197.

The eluant can be any suitable eluant that can displace the adsorbedprecious metal from the loaded resin beads. The eluant could includesalts, such as one or more types of polythionate ions as set forth inU.S. application Ser. No. 09/542,736 above, and a nitrate, athiocyanate, a sulfite, a thiourea, a perchlorate and mixtures thereof.

Typically, the concentration of the eluant in the pregnant solution 196ranges from about 0.25 to about 3 M; the temperature of elution 194 fromabout 5 to about 70° C., and the pH of elution 194 from about pH 5 toabout pH 12. Under the conditions, at least about 90% and more typicallyfrom about 95 to about 99% of the precious metal adsorbed on the resinis displaced by the eluant into the pregnant solution 196.

The stripped resin 197 is recycled to the resin-in-pulp step 164.Optionally, the stripped resin 197 can be regenerated (not shown) byknown techniques prior to reuse of the resin. As will be appreciated,the resin can be regenerated by acid washing the resin with an acid suchas nitric acid or hydrochloric acid. The acid wash removes adsorbedeluant and/or impurities from the resin and frees up the functionalsites on the resin surface (previously occupied by the eluant) to adsorbadditional precious metal. In the case of a polythionate eluant, theresin can be regenerated by contacting the resin with sulfide and/orsulfite to reduce the polythionate ions to thiosulfate ions and sulfateions. After regeneration, the resin and regeneration product solutionare separated by screening and washing.

The pregnant solution 196, which includes the eluant and typically nomore than about 100 ppm and more typically from about 10 to about 500ppm solubilized precious metals, is subjected to electrowinning 198 torecover the solubilized precious metals and form a barren solution 199.Problems in electrowinning of precious metals out of a medium containingpolythionates and/or thiosulfate have been encountered in U.S. patentapplication Ser. No. 09/542,736. When the eluant is a polythionate thepolythionate and thiosulfate tend to be co-reduced with the preciousmetal at the cathode to produce elemental sulfur, which interferes withthe efficient continued operation of the electrowinning circuit whilethe polythionate and thiosulfate are also wastefully oxidized to sulfateions at the anode.

These problems are overcome by the present invention through the use ofsulfite in the pregnant solution. Sulfite is added to the eluant and/orto the pregnant solution 196 prior to, during, or after electrowinning.Preferably, sulfite is added to the eluant prior to the elution step194. In the presence of sulfite, the precious metal is reduced at thecathode while the sulfite is oxidized to sulfate at the anode. This hasthe benefit of lowering the cell voltage required. Preferably, theconcentration of sulfite in the pregnant solution 196 (in the elutionand electrowinning steps 194, 198) is at least about 0.01M and morepreferably ranges from about 0.1 to about 2 M. The sulfite is preferablyin the pregnant solution with another eluant, such as any of the eluantsnoted above.

The stripped or barren solution 199 is removed from the electrowinningcell(s) and returned to the elution step 194. A bleedstream (not shown)of the barren solution 199 can be used to control buildup of impuritiessuch as sulfate.

The recovered precious metal 195, which contains the precious metalrecovered in electrowinning and impurities, is subjected to retorting193 by known techniques to remove the impurities and form precious metalsludge. The sludge, which contains at least most of the precious metalin the recovered precious metal 195, is refined to produce a preciousmetal product of high purity.

FIG. 2 depicts another embodiment of a process for thiosulfate leachingof a refractory precious metal-containing material. FIG. 2 shows analternative to resin-in-pulp for precious metal recovery. Followingleaching 132, the precious metal bearing solution 144 is separated 200from the solids by any suitable means, such as by counter-currentdecantation washing and/or filtration. Preferably, at least about 95%and more preferably at least about 99% of the precious metal isseparated from the solids with the latter going to tailings impoundment.

The separated precious metal bearing solution 204 is directed to theprecious metal precipitation—thiosulfate regeneration step 208. Thisprocess can be carried out in any suitably agitated reactor or pluralityof agitated reactors. The pH of the precious metal bearing solution 204is adjusted if necessary to about pH 5.5-12, more preferably about pH7-11, even more preferably about pH 9-11 using a suitable basic reagentsuch as sodium hydroxide and the solution is contacted with a reductant,preferably a sulfide and/or bisulfide and/or polysulfide reagent toprecipitate at least about 99% of the precious metal and convert atleast about 90% of the polythionates to thiosulfate, effectivelyregenerating the thiosulfate lixiviant. The effectiveness of theconversion causes significantly less thiosulfate reagent to be consumedduring the process than for conventional thiosulfate leaching processes.The use of a sulfide and/or bisulfide and/or polysulfide has the addedbenefit of reducing impurities such as copper or mercury or manganesefrom solution thereby reducing the rate of thiosulfate degradation.While not wishing to be bound by any theory, it is believed that themost likely composition of the precipitate is the metallic preciousmetal and/or a precious metal sulfide, such as Au₂S. Maximumprecipitation of gold and regeneration of thiosulfate is accomplished byadding at least a stoichiometric amount of reductant (relative to thedissolved precious metal and polythionate concentrations) to reduce thesolution ORP to at least about −150 mV (SHE). The temperature ispreferably maintained in the range of about 5 to 40° C., and morepreferably at ambient temperature, about 20° C. The retention time isabout 5 minutes to about 2 hours, more preferably about 15 minutes toabout 1 hour. The process is conducted under oxygen-depleted conditions,with the solution preferably containing no more than about 1 ppmdissolved molecular oxygen and more preferably less than about 0.2 ppmdissolved molecular oxygen concentration, by bubbling anoxygen-deficient gas such as nitrogen into the slurry and/or maintaininga blanket of nitrogen in the atmosphere over the slurry as noted above.

The precious metal bearing precipitate is separated from the regeneratedsolution 212 by any suitable method such as filtration, CCD, and thelike and the separated precious metal 216 is recovered by refining infurnaces.

The regenerated solution 220 is directed to the conditioning step 224,which can be conducted in any suitably agitated reactor or plurality ofreactors. The solution pH is adjusted to a value suitable for recyclingthe solution back to grinding 108 and/or for precious metal scavenging228. Preferably, the pH ranges from about pH 7 to about pH 12, morepreferably about pH 8 to pH 10. The solution 220 is agitated in thepresence of an oxygen-containing atmosphere, such as air, to oxidize anyremaining reductant (such as sulfide or bisulfide or polysulfide)carried over from the precious metal precipitation—thiosulfateregeneration step 208. The duration of the conditioning step 224 ispreferably not sufficient to cause more than about 5% of the thiosulfateto form polythionates, or to yield a polythionate concentration of morethan about 0.003M. The majority (typically at least about 80 vol %) ofthe conditioned solution 232 is then recycled in recycle solution 236. Aminor portion (e.g., from about 2 to about 20 vol %) of the conditionedsolution or bleed stream 240 may have to be bled to tailings to controlthe buildup of impurities, such as soluble sulfate and metallicimpurities. Prior to discharge to tailings the bleed portion 240 of theconditioned solution 232 is directed to the precious metal scavengingstep 228 to recover any precious metals remaining in solution that werenot recovered in the precious metal precipitation—thiosulfateregeneration step 208. Precious metal scavenging can be accomplished, byany suitable gold recovery technique such as by passing the bleedsolution 240 through a column containing a strong base resin to adsorbthe precious metal. While not wishing to be bound by any theory,precipitated precious metal can be redissolved due to trace amount ofmolecular oxygen in the solution and incomplete reduction ofpolythionates in the solution. Because the amount of polythionates inthe bleed is negligible, a resin-in-column recovery technique will havean excellent ability to load any remaining dissolved precious metal.

In an alternative configuration (not shown), the precious metalprecipitates are redissolved in a suitable solvent, such asnitric/hydrochloric acid, cyanide, thiosulfate, thioureachloride/chlorine and bromide/bromine to provide a preciousmetal-containing solution. The precious metal can then be recovered byelectrolysis as noted above in connection with step 198 of FIG. 1.

This process is preferred in certain applications over the process ofFIG. 1. For certain precious metal-containing materials, it is difficultto obtain high rates of precious metal adsorption onto resins whilemaintaining the precious metal in solution. The use of an RIPpretreatment step, though beneficial, can be difficult to use withoutexperiencing some precious metal precipitation. Conditioning 156 may notbe completely effective in redissolving gold precipitates, which wouldbe discarded with the barren solids to tailings. The process of FIG. 2can also be more robust, simpler, and therefore easier to design andoperate than the process of FIG. 1.

FIG. 3 shows an alternative to FIG. 2 in which thiosulfate leaching isconducted in two stages to achieve more effective recovery of theprecious metal content. Leaching is first conducted at atmosphericpressure and ambient temperature in the presence of an oxygen-containinggas such as air or industrially available oxygen (step 300) to dissolvefrom about 30 to 95% of the leachable precious metal content. Theleachable precious metal content is defined as that portion of theprecious metal content that is physically accessible to the thiosulfatelixiviant and is not encapsulated within constituents contained in thehost material. The precious metal bearing solution 304 is separated fromthe solids 308 (step 200), the solids 308 are repulped with a portion310 of the recycle solution 236, and the resulting slurry 308 is thendirected to pressure leaching (step 312) to dissolve the majority, ie.about 5-70%, of the remaining leachable precious metal content that wasnot recovered in atmospheric leaching 300. In pressure leaching thesolids are leached under superatmospheric conditions such as theconditions described previously (step 132 of FIG. 1). The molecularoxygen partial pressure in leach 300 preferably ranges from themolecular oxygen partial pressure at ambient conditions (e.g., more thanabout 3 psia at sea level) to about 15 psia and the molecular oxygenpartial pressure in leach 312 preferably ranges from more than 15 psiato about 500 psia. The slurry 316 exiting pressure leaching 312 is thenprocessed in essentially the same manner as the slurry exiting leaching300 in FIG. 2. That is, the slurry 316 is subjected to solid/liquidseparation 320 in the presence of wash water to separate the barrensolid material 324 from the (second) pregnant leach solution 328. Thefirst and second pregnant leach solutions 304, 328 are subjected toprecious metal precipitation—thiosulfate regeneration 208, furthersolid/liquid separation 212, conditioning 224 and precious metalscavenging 228 as noted above in connection with FIG. 2.

The process of FIG. 3 typically performs the bulk of the leaching, orprecious metal dissolution, under ambient conditions, which is muchcheaper than leaching under superatmospheric conditions. Themore-difficult-to-dissolve precious metals and weakly preg-robbedprecious metals are then dissolved in a higher pressure leach. Becauseless precious metal remains to be dissolved, the high pressure leach canhave a shorter residence time and therefore lower capacity than would bepossible in the absence of the ambient pressure leach.

FIG. 4 depicts another embodiment of the present invention. The processis similar to those discussed above except that thiosulfate leaching isperformed by heap leaching 400 techniques. The comminuted preciousmetal-containing material 404 can be directly formed into a heap (inwhich case the material would have a preferred P₈₀ size of from about 2inches to about ¼ inch, possibly agglomerated and formed into a heap.The thiosulfate lixiviant (which commonly includes a recycledthiosulfate lixiviant 236 mixed with a makeup (fresh) thiosulfatesolution (not shown)) is applied to the top of the heap usingconventional techniques, and the pregnant leach solution 408 iscollected from the base of the heap. Refining can be performed using anyof the techniques noted above.

To facilitate extraction of gold from sulfidic and/or carbonaceousmaterials, the thiosulfate leach step in any of the above processes canbe preceded by one or more pretreatment steps to destroy or neutralizethe carbon-containing and/or sulfidic minerals. By way of example, theintermediate steps can include one or more of biooxidation or chemicaloxidation to oxidize sulfides, ultrafine grinding to liberate occludedprecious metals, conventional roasting to destroy carbon- and/orsulfide-containing minerals, and/or microwave roasting.

Example 1

A gold ore from Nevada, designated Sample A, was subjected tothiosulfate leaching under oxygen pressure at varying temperatures. Theore assayed 24.1 g/t gold, 2.59% iron, 0.31% total sulfur, 0.28% sulfidesulfur, 3.40% total carbon, 1.33% organic carbon and 0.02% graphiticcarbon. From a diagnostic leaching analysis of the ore it was determinedthat a maximum of 83% of the contained gold was capable of beingsolubilized while the remaining gold was inaccessible to a lixiviantbecause it was encapsulated within pyrite and/or other mineralscontained in the ore.

The ore was ground to 80% passing 200 mesh (75 μm). Samples of the orewere slurried with water to a pulp density of 33% solids in amechanically agitated laboratory autoclave. The natural pH of the slurryat ambient temperature was 8.3. The pH of the slurry was adjusted to 9with sodium hydroxide and a quantity of sodium thiosulfate reagent wasadded to adjust the initial leach solution thiosulfate concentration to0.1 molar (M). The autoclave was sealed and pressurized to 100 psigoxygen with pure (95% plus) oxygen gas and the slurry was heated to thedesired temperature (if required). Leaching was maintained for 6 hours,during which pulp samples were taken at 2 and 4 hours in order tomonitor gold extraction with time. Upon termination of leaching, theslurry was filtered and the residue solids were washed with a dilutethiosulfate solution. The residue solids and leach solution were assayedfor gold to determine the final gold extraction.

The results were as follows:

Leach Temp. Leach Time Calc'd Head Residue Au Ext'n (° C.) (hours) Au(g/t) Au (g/t) (%) 20 2 33.3 4 41.9 6 22.8 9.44 58.5 40 2 51.2 4 55.1 626.4 9.25 64.9 60 2 63.7 4 68.5 6 22.8 4.26 81.3 60 (repeat) 2 65.2 473.0 6 80.9

The results indicate that the rate and extent of gold extraction wasimproved with increasing temperature and leach time in the temperaturerange 20-60° C. The best results were obtained at 60° C., with about 81%gold extraction obtained after 6 hours leaching, this representing about98% of the leachable gold content of the ore.

Example 2

A second gold ore from Nevada, designated Sample B, was subjected tothiosulfate leaching under oxygen pressure at varying initial pH's. Theore assayed 9.45 g/t gold, 2.50% iron, 0.39% total sulfur, 0.36% sulfidesulfur, 4.20% total carbon, 1.46% organic carbon and 0.05% graphiticcarbon. From a diagnostic leaching analysis of the ore it was determinedthat 82% of the contained gold was capable of being solubilized. Samplesof the ore were prepared and leached as described in Example 1, exceptthe temperature was 60° C. in each test, the autoclave was pressurizedwith 50 psig oxygen, the initial pH was adjusted to either 9, 11 or 12,and the leach retention time was extended to 8 hours for the pH 11 and12 tests.

The results were as follows:

Initial Leach Time Calc'd Head Residue Au Ext'n pH (hours) Au (g/t) Au(g/t) (%) 9 1 50.2 2 62.4 4 72.0 6 8.49 2.10 75.3 11 1 41.3 2 63.0 469.3 8 8.61 2.00 76.8 12 1 6.4 2 1.0 4 13.6 8 8.61 3.34 61.2

The results indicate that there was not much difference in gold leachingbehaviour over the initial pH range of 9-11 (it should be noted that thepH tended to decline during leaching). However, gold leaching wassuppressed during the first 4 hours of leaching at pH 12, but thenstarted to recover.

Example 3

A third gold ore sample from Nevada, Sample C, was subjected tothiosulfate leaching under oxygen pressure at varying temperatures. Thehead analysis of the ore was as follows:

Gold Ore Sample C Au, g/t 9.50 C (t), % 4.45 Fe, % 2.52 C (CO₃), % 3.12Cu, ppm 40 C (org), % 1.38 As, ppm 647 S (2-), % 0.35 Hg, ppm 14 S (t),% 0.27 Ca, % 9.0 Mg, % 1.5From a diagnostic leaching analysis of the ore it was determined that83% of the contained gold was capable of being solubilized.

The ore was ground to 80% passing 200 mesh (75 μm). Samples of the orewere slurried with water to a pulp density of 33% solids in amechanically agitated laboratory autoclave. The initial pH of the slurrywas adjusted to approximately 11 with sodium hydroxide, after which theautoclave was sealed and pressurized to 100 psig oxygen with pure (95%plus) oxygen gas and the slurry was heated to the desired temperature.To initiate leaching, a quantity of sodium thiosulfate stock solutionwas injected to adjust the leach solution thiosulfate concentration to0.1 M. Leaching was continued for 6 to 10 hours, during which noadditional reagents were added. Pulp samples were taken at set intervalsduring leaching in order to monitor gold extraction with time. Upontermination of leaching, the slurry was filtered and the residue solidswere washed with a dilute thiosulfate solution. The residue solids andleach solution were assayed for gold to determine the final goldextraction.

FIG. 5 depicts graphically the effect of leach temperature, in the range40-80° C., on the rate of gold extraction from Sample C. It can be seenthat the gold leached quickly at 60° C. and 80° C., there being littledifference in the extraction rate at the two temperatures. The goldextraction peaked at approximately 83%, the maximum extractable, after 6hours leaching. Gold leaching was slowed if the temperature was loweredto 40° C., but 80% gold extraction was still obtained after 10 hoursleaching at 40° C.

An overall summary of the results is provided below:

Test #6 Test #25 Test #15 Parameter 80° C. 60° C. 40° C. Leach time,hours 8 6 10 Final pH 7.0 8.7 9.3 Final ORP, mV (SHE) 307 242 225 Calc'dHead Au, g/t 9.48 9.43 9.27 Residue Au, g/t 1.59 1.63 1.81 Au Ext'n, %83.2 82.7 80.5

Example 4

The gold ore designated Sample C was subjected to thiosulfate leachingat varying oxygen pressures. Samples of the ore were prepared andleached as described in Example 3 except the temperature was maintainedat 60° C. in each test and the oxygen partial pressure was varied.

FIG. 6 portrays the effect of oxygen partial pressure, in the range0-200 psig, on the rate of gold extraction from Sample C (in the 0 psigO₂ test, the autoclave was not pressurized but the head space wasmaintained with pure oxygen at atmospheric pressure). It can be seenthat the rate of gold leaching was somewhat sensitive to oxygenpressure, in that the rate increased with increasing pressure,particularly during the first two hours of leaching. After 6 hoursleaching, gold extraction varied from a low of 78% at 0 psig O₂ to ahigh of 83% at 200 psig O₂.

An overall summary of the results is provided below:

Test #7 Test #25 Test #22 Test #28 Test #31 Parameter 200 psig O₂ 100psig O₂ 50 psig O₂ 10 psig O₂ 0 psig O₂ Leach time, hours 8 6 6 6 6Final pH NA 8.7 9.0 9.3 9.4 Final ORP, mV (SHE) NA 242 223 216 232Calc'd Head Au, g/t 9.78 9.43 9.40 8.95 9.08 Residue Au, g/t 1.68 1.631.77 1.72 2.00 Au Ext'n, % 82.8 82.7 81.1 80.8 78.0 NA = not analyzed

Example 5

The gold ore designated Sample C was subjected to thiosulfate leachingunder oxygen pressure at varying initial sodium thiosulfateconcentrations. Samples of the ore were prepared and leached asdescribed in Example 3 except the temperature was maintained at 60° C.in each test and the initial sodium thiosulfate concentration wasvaried.

FIG. 7 portrays the effect of initial sodium thiosulfate concentration,in the range 0.05-0.2 M, on the rate of gold extraction from Sample C.It can be seen that the rate of gold leaching was insensitive to initialthiosulfate concentration in the 0.1-0.2 M range. At 0.05 M thiosulfate,the rate was reduced significantly, particularly during the first twohours of leaching. After 6 hours leaching gold extraction was 78% at0.05 M thiosulfate compared to 82% achieved at both 0.1 M and 0.2 Mthiosulfate concentration.

An overall summary of the results is provided below:

Test #4 Test #25 Test #8 Parameter 0.2M 0.1M 0.05M Leach time, hours 8 66 Final pH 8.7 8.7 8.5 Final ORP, mV (SHE) NA 242 262 Calc'd Head Au,g/t 8.85 9.43 9.40 Residue Au, g/t 1.50 1.63 1.87 Au Ext'n, % 83.0 82.780.1 NA = not analysed

Example 6

The gold ore designated Sample C was subjected to thiosulfate leachingunder oxygen pressure at two different pulp densities. Samples of theore were prepared and leached as described in Example 3, except thetemperature was maintained at 60° C. in each test and the leach pulpdensity was either 33% or 45% solids by weight.

FIG. 8 portrays the effect of 33% vs. 45% pulp density on the rate ofgold extraction from Sample C. The rate of gold leaching was found to beessentially insensitive to pulp density in this range.

An overall summary of the results is provided below:

Test #26 Test #25 45% pulp 33% pulp Parameter density density Leachtime, hours 6 6 Final pH 8.5 8.7 Final ORP, mV (SHE) 286 242 Calc'd HeadAu, g/t 9.29 9.43 Residue Au, g/t 1.68 1.63 Au Ext'n, % 81.9 82.7

Example 7

A fourth gold ore sample from Nevada, Sample D, was subjected tothiosulfate leaching at 60° C. and 10 psig oxygen pressure at thenatural pH of the slurry, for 8 hours. The head analysis of the ore wasas follows:

Gold Ore Sample D Au, g/t 12.15 C (t), % 4.31 Fe, % 2.09 C (CO₃), % 3.02Cu, ppm 39 C (org), % 1.30 As, ppm 692 S (2-), % 0.12 Hg, ppm 27 S (t),% 0.22 Ca, % 8.9 Mg, % 1.3From a diagnostic leaching analysis of the ore it was determined that80% of the contained gold was capable of being solubilized.

The ore was ground to 80% passing 200 mesh (75 μm). A sample of the orewas slurried with water to a pulp density of 40% solids in amechanically agitated laboratory autoclave. The autoclave was sealed andpressurized to 100 psig oxygen with pure (95% plus) oxygen gas and theslurry was heated to 60° C. To initiate leaching, a quantity of sodiumthiosulfate stock solution was injected to adjust the leach solutionthiosulfate concentration to 0.1 M. Leaching was continued for 8 hours,during which no additional reagents were added. Pulp samples were takenat set intervals during leaching in order to monitor gold extraction andremaining thiosulfate with time. Upon termination of leaching, theslurry was filtered and the residue solids were washed with a dilutethiosulfate solution. The residue solids and leach solution were assayedfor gold to determine the final gold extraction.

FIG. 9 depicts percent gold extraction and percent remaining thiosulfatewith time. Gold extraction reached 79.3% after 8 hours, or 99% of theleachable gold content. Thiosulfate consumption was low, with 86.7% ofthe thiosulfate remaining after 8 hours and available for reuse.

An overall summary of the results is provided below:

Parameter Test #37-01 Leach time, hours 8 Initial pH 7.9 Final pH 9.0Initial ORP, mV (SHE) 411 Final ORP, mV (SHE) 397 Calc'd head Au, g/t11.50 Residue Au, g/t 2.38 Gold extraction, % 79.3 Amount of thiosulfateremaining, % 86.7

Example 8

A thiosulfate leach discharge slurry was heated to 60° C. in an agitatedreactor in preparation for RIP pre-treatment, the objective being toreduce the polythionate content without precipitating gold. The slurrywas kept under a nitrogen atmosphere to ensure the dissolved oxygencontent was maintained below 0.2 mg/L. A single dose of a 0.26 M sodiumbisulfide (NaHS) solution, adjusted to pH 9, was added and thepretreatment was allowed to proceed at 60° C. and ambient pressure for 2hours. The amount of sulfide added was 150% of stoichiometric based onthe amount required to convert the polythionates back to thiosulfate inaccordance with the following reactions:

2S₄O₆ ²⁻+S²⁻+3/2H₂O→9/2S₂O₃ ²⁻+3H⁺

S₃O₆ ²⁻+S²⁻→2S₂O₃ ²⁻

A summary of the results is provided below:

Time Au S₂O₃ ²⁻ S₄O₆ ²⁻ S₃O₆ ²⁻ ORP (min) (mg/L) (g/L) (g/L) (g/L) (mV)pH 0 4.36 8.38 0.51 0.59 240 6.9 120 4.36 11.0 0.06 0.10 5 6.7

The tetrathionate and trithionate concentrations were reduced by 88% and83% respectively while all of the gold remained in solution.

Example 9

A pregnant thiosulfate leach solution was adjusted to pH 10 with sodiumhydroxide in an agitated reactor in preparation for sulfide treatment,the objective being to regenerate thiosulfate and precipitate the gold.The solution was kept under a nitrogen atmosphere to ensure thedissolved oxygen content was maintained below 0.2 mg/L. A single dose ofa 0.26 M sodium sulfide (Na₂S) solution was added and the treatment wasallowed to proceed for 2 hours at ambient temperature (22° C.) andpressure. The amount of sulfide added was 100% of stoichiometric basedon the amount required to convert the polythionates back to thiosulfatein accordance with the following reactions:

2S₄O₆ ²⁻+S²⁻+3/2H₂O→9/2S₂O₃ ²⁻+3H⁺

S₃O₆ ²⁻+S²⁻→2S₂O₃ ²⁻

A summary of the results is provided below:

Time Au S₂O₃ ²⁻ S₄O₆ ²⁻ S₃O₆ ²⁻ ORP (min) (mg/L) (g/L) (g/L) (g/L) (mV)pH 0 4.12 7.8 0.84 1.47 200 10.0 10 0.05 9.9 0.01 0.01 −210 11.0 20 0.029.9 0.01 0.01 −220 10.4 30 0.01 9.9 0.01 0.01 −230 10.2 60 0.01 9.8 0.010.01 −260 10.3 90 0.01 10.1 0.01 0.01 −260 10.3 120 0.01 10.2 0.01 0.01−260 10.3

The rate of conversion of polythionates to thiosulfate was extremelyfast under ambient conditions, with essentially complete conversionachieved after about 10 minutes. Similarly, gold precipitation was alsofast and essentially complete after about 30 minutes.

While this invention has been described in conjunction with the specificembodiments thereof, it is evident that many alternatives,modifications, and variations will be apparent to those skilled in theart. Accordingly, preferred embodiments of the invention as set forthherein are intended to be illustrative, not limiting. By way of example,any source of sulfur species with an oxidation state less than +2 may beused in any of the above process steps to convert polythionates tothiosulfate. The regeneration step 184 in FIG. 1 can be performed in avariety of locations. For example, regeneration 184 can be performed inthe recycle loop after CCD 172 and before grinding 108, between grinding108 and thickening 116, in the thickener 116 immediately before orduring, leaching 132 and/or between resin in pulp 164 and CCD 172. Freshthiosulfate can also be added in a number of locations. For example,fresh thiosulfate can be added in any of the locations referred topreviously for the regeneration step 184 and can be added after orduring regeneration 184 as noted above or in a separate tank orlocation. In FIG. 3, a lixiviant other than thiosulfate, such ascyanide, can be used in the atmospheric leach 300 with thiosulfate beingused in the pressure leach 312. These and other changes may be madewithout departing from the spirit and scope of the present invention.

1-106. (canceled)
 107. A method, comprising: (a) contacting a preciousmetal-containing material with a thiosulfate lixiviant to dissolve theprecious metal and form a pregnant thiosulfate leach solution containingthe dissolved precious metal and a polythionate; (b) contacting thepregnant thiosulfate leach solution with at least one of asulfite-containing reagent and a sulfide-containing reagent to convertat least most of the polythionates into thiosulfate; (c) thereaftercontacting the pregnant leach solution with an extraction agent tocollect at least most of the dissolved precious metal on the extractionagent and form a barren thiosulfate leach solution for recycle to step(a); and (d) recovering the collected precious metal from the preciousmetal loaded extraction agent.
 108. The method of claim 107, furthercomprising, after step (b) and before step (c): conditioning thepregnant leach solution in an oxidizing atmosphere to redissolve anyprecipitated precious metal sulfides.
 109. The method of claim 107,wherein polythionates compete with dissolved precious metal complexesfor loading on the extraction agent.
 110. The method of claim 108,wherein the at least one of a sulfite-containing reductant andsulfide-containing reductant is the sulfite-containing reductant andwherein the sulfite-containing reductant is at least one of ametabisulfite and sulfur dioxide.
 111. The method of claim 108, whereinthe at least one of a sulfite-containing reductant andsulfide-containing reductant is the sulfide-containing reductant andwherein the sulfide-containing reductant is at least one of a bisulfide,sulfide, hydrogen sulfide gas, and a polysulfide.
 112. A method,comprising: (a) contacting a precious metal-containing material with athiosulfate lixiviant to dissolve the precious metal and form a pregnantthiosulfate leach solution containing the dissolved precious metal; (b)recovering the dissolved precious metal from the pregnant thiosulfateleach solution to form a barren leach solution; and (c) sparging an atleast substantially non-oxidizing gas through at least one of thepregnant thiosulfate leach solution and barren leach solution to inhibitthe formation of polythionates.
 113. The method of claim 112, whereinthe substantially non-oxidizing gas is substantially free of oxidants.114. The method of claim 112, wherein the substantially non-oxidizinggas comprises at least about 85 vol. % of an inert gas and no more thanabout 5 vol. % oxidants. 115-134. (canceled)
 135. The method of claim114, wherein the inert gas in the substantially non-oxidizing gas isnitrogen gas.
 136. The method of claim 112, wherein the substantiallynon-oxidizing gas comprises a reducing gas.
 137. The method of claim113, wherein the substantially non-oxidizing gas comprises no more thanabout 5 vol. % oxidants.